JA Filed by Filing Services Canada Inc. 403-717-3898

1.1	Summary	16
1.1	Introduction	16
1.2	Terms of Reference	16
1.3	Project Setting	16
1.4	Mineral Tenure, Royalties and Agreements	18
1.5	Geology and Mineralization	18
1.6	History	18
1.7	Drilling and Sampling	18
1.8	Mineral Resource Estimate	19
1.9	Mineral Resource Statement	19
1.10	Mining Methods	21
1.11	Metallurgy and Recovery Methods	21
1.12	Project Infrastructure	21
1.13	Markets and Contracts	24
1.14	Environmental, Permitting and Social Considerations	24
1.15	Capital Cost Estimates	25
1.16	Operating Cost Estimates	25
1.17	Economic Analysis	26
1.18	Interpretations and Conclusions	27
2	Introduction	28
2.1	Terms of Reference	28
2.2	Qualified Persons	29
2.3	Site Visits and Scope of Personal Inspections	29
3	Reliance on Other Experts	31
4	Property Description and Location	32
5	Accessibility, Climate, Local Resources, Infrastructure and Physiography	36
6	History	37
6.1	Historical Resource Estimates	39
7	Geological Setting and Mineralization	43
7.1	Regional Geology	43
7.2	Property Geology	43
7.3	Mineralization	47
8	Deposit Types	49
9	Exploration	50
10	Drilling	52
11	Sample Preparation, Analyses and Security	70
11.1	Sample Method and Approach	70
11.2	2006 Program Sampling	71
12	Data Verification	80
13	Mineral Processing and Metallurgical Testing	81
13.1	Introduction	81

13.2	Metallurgical Test Work History	81
13.3	Samples	83
13.4	Mineralogy	85
13.5	RDI 2003	85
13.6	Mintek 2003	90
13.7	RDI 2004	91
13.8	Oxidor 2004	92
13.9	PRA 2007	93
13.10	Outotec 2018	101
13.11	Recommended Flowsheet	106
13.12	Metallurgical Performance Projections	106
14	Mineral Resource Estimates	109
14.1	Introduction	109
14.2	Resource	109
14.3	Database and Assumptions of the Resource Estimate	110
14.4	Assay Statistics, Capping and Compositing	113
14.5	Variography	116
14.6	Block Model	116
14.7	Block Model Validation	120
15	Mineral Reserve Estimates	126
16	Mining Method	127
16.1	Introduction	127
16.2	Mine Planning Criteria	127
16.3	Geotechnical Criteria	128
16.4	Mine Production Plan	128
16.5	Mining Methods	130
16.6	Mine Design	134
16.7	Mine Ventilation	135
16.8	Underground Mine Services	137
16.9	Mine Safety	138
16.10	Paste Backfill	139
16.11	Mine Equipment	139
16.12	Mine Personnel	140
17	Recovery Methods	142
17.1	Process Flowsheet	142
17.2	Process Design Criteria	142
17.3	Process Description	145
17.4	Reagents and Power Consumption	147
17.5	Process Water and Power	148
18	Project Infrastructure	149
18.1	Overall Site Development	149
18.2	Site Services	151
18.3	Mobile Fleet	153
18.4	Buildings and Facilities	154

19	Market Studies and Contracts	158
20	Environmental Studies, Permitting and Social or Community Impact	159
20.1	Environmental	159
20.2	Regulatory Framework	160
20.3	Socio-Economic	165
21	Capital and Operating Costs	167
21.1	Capital Cost Estimate	167
21.2	Operating Cost Estimate	173
22	Economic Analysis	178
22.1	Economic Sensitivity	180
22.2	Cautionary Statement	181
23	Adjacent Properties	183
24	Other Relevant Data and Information	184
25	Interpretation and Conclusions	185
25.1	Mineral Tenure, Surface Rights, Water Rights, Royalties and Agreements	185
25.2	Geology, Mineralization, Exploration	185
25.3	Mineral Resource	185
25.4	Mining	186
25.5	Metallurgy and Process	186
25.6	Environmental, Permitting and Social Impact	187
25.7	Infrastructure and Costing	187
25.8	Costs and Economic Analysis	187
26	Recommendations	189
26.1	Resource Recommendations	190
26.2	Mining Recommendations	190
26.3	Process Recommendations	190
26.4	Infrastructure Recommendations	190
26.5	Environmental, Permitting and Social Impact	191
27	References	192

LIST OF TABLES

Table 1-1	New Polaris Indicated Resources	20
Table 1-2	New Polaris Inferred Resources	20
Table 1-3	Capital Cost Summary	25
Table 1-4	Operating Cost Summary	25
Table 1-5	PEA Summary	27

Table 4-1	List of Claims	33

Table 6-1	Summary of Exploration Drilling to 2006	38
Table 6-2	Historic Resource Estimates, 1988 to 1995	41
Table 6-3	Historic Resource Estimate, 2007	42

Table 9-1	Exploration Employees / Contractors	51

Table 10-1	Diamond Drillholes (1988 to 2005)	53
Table 10-2	Assay Composites C Vein System	60

Table 11-1	Univariate Statistics, Blank Samples	71
Table 11-2	Univariate Statistics, Duplicate Samples	72
Table 11-3	Standard Samples	74
Table 11-4	Univariate Statistics, Round Robin Samples	77

Table 13-1	New Polaris Metallurgical Test Work History	81
Table 13-2	RDI Bulk Head Grade	83
Table 13-3	RDI Bulk Sample Mineralogical Data – XRD Results	84
Table 13-4	PRA Sample Grades	84
Table 13-5	PRA Sample Mineralogical Data	84
Table 13-6	Inspectorate Sample (Canarc 2015)	84
Table 13-7	Summary of Rougher Flotation Tests – Adapted from RDI 2003	87
Table 13-8	Summary of Cleaner Flotation Tests – Adapted from RDI 2003	88
Table 13-9	Summary of Gravity Test (Rougher Tails) – Adapted from RDI 2003	89
Table 13-10	Summary of Gravity Test (Rougher Conc.) – Adapted from RDI 2003	89
Table 13-11	Leach Test Results – Adapted from RDI 2003	89
Table 13-12	Rougher Concentrate Production Analysis – Adapted from RDI 2003	90
Table 13-13	Results of Diagnostic Leach – Adapted from Mintek 2003	91
Table 13-14	Results of Batch Acid POX Tests – Adapted from RDI 2004	92
Table 13-15	Gravity Concentrate Test Results	93
Table 13-16	Gravity Tails Bulk Flotation Test Results – Adapted from PRA 2007	94
Table 13-17	As/Sb Depression Flotation Test Results – Adapted from PRA 2007	95
Table 13-18	Rougher Regrind Cleaner Flotation Test Results – Adapted from PRA 2007	96
Table 13-19	Test F8 Flotation Test Results – Adapted from PRA 2007	97

Table 13-20	Locked Cycle Flotation Test Results – Adapted from PRA 2007	100
Table 13-21	Summary of BIOX™ BAT and Bottle Roll Tests – Adapted from Outotec 2018	102
Table 13-22	Summary of Neutralization Tests – Adapted from Outotec 2018	103
Table 13-23	Summary of Flocculant Screening Tests – Outotec 2018	104
Table 13-24	Summary of Flocculant Screening Tests – Outotec 2018	105
Table 13-25	New Polaris Projected Metallurgical Recoveries	107

Table 14-1	Summary of Indicated and Inferred Total Resource	110
Table 14-2	Summary of Assay Data in the Project Area	111
Table 14-3	Summary of Assay Data within the Domains used for Resource Estimation	111
Table 14-4	Summary of Capping Value by Vein for Gold	113
Table 14-5	Summary of Assay and Composite Statistics	115
Table 14-6	Variogram Parameters	116
Table 14-7	Block Model Parameters	117
Table 14-8	Summary of Search Orientation and Distance for each Interpolation Pass	118
Table 14-9	Sample Selection Criteria during Interpolation	118
Table 14-10	Classification Criteria	119
Table 14-11	Summary of Model and De-Clustered Composite (NN) Mean Grades	121

Table 16-1	Mine Planning Criteria	128
Table 16-2	New Polaris Mine Production and Development Schedule	129
Table 16-3	Diesel Equipment Ventilation Requirements	136
Table 16-4	Mine Personnel	140

Table 17-1	Summary of New Polaris Preliminary Process Design Criteria	144
Table 17-2	Summary of Reagents and Consumables	148

Table 18-1	Total Site Power Requirements	151
Table 18-2	Estimated Annual Project Fuel Requirements	152
Table 18-3	Plant Mobile Equipment Fleet	153
Table 18-4	Surface Mobile Equipment Fleet	154

Table 21-1	Project Capital Cost Estimate	167
Table 21-2	Overall Site Capital Cost Estimate	168
Table 21-3	Mining Capital Cost Estimate	168
Table 21-4	Process Plant Capital Cost Estimate	170
Table 21-5	CDF Capital Cost Estimate	170
Table 21-6	Site Services Capital Cost Estimate	171
Table 21-7	Buildings and Facilities Capital Cost Estimate	171
Table 21-8	Indirect Capital Cost Estimate	172
Table 21-9	Project Operating Cost Estimate	174
Table 21-10	Mine Operating Cost Estimates	174
Table 21-11	Process Operating Cost Estimate	175
Table 21-12	G&A Cost Estimate	175

LIST OF FIGURES

Figure 1-1	Location Map	17
Figure 1-2	New Polaris General Arrangement	22
Figure 1-3	New Polaris General Arrangement Map with Claim Boundaries	23
Figure 4-1	Claim Location Map	34
Figure 4-2	Claim Map with Principle Vein Locations	35
Figure 7-1	Regional Geology	45
Figure 7-2	Property Geology	46
Figure 7-3	3D Model C Vein; plan view in mine specific space	48
Figure 10-1	Drillhole Locations, Plan Map	64
Figure 10-2	Drillhole Locations and Modelled 3D solids; Sectional View in mine specific space, looking east	65
Figure 10-3	Drillhole Locations and Modelled 3D solids; Sectional View in mine specific space, looking north	66
Figure 10-4	Drillhole Locations and Modelled 3D solids, North South Cross-section, East 1766 in mine specific space, looking west	67
Figure 10-5	Drillhole Locations and Modelled 3D solids, North South Cross-section, East 1916 in mine specific space, looking west	68
Figure 10-6	Drillhole Locations and Modelled 3D solids , West East Cross-section, North 867.5 in mine specific space, looking north	69
Figure 11-1	Blank Samples, 2006	72
Figure 11-2	Duplicate Samples, 2006	73
Figure 11-3	Standard PM 165, 2006	74
Figure 11-4	Standard PM 415, 2006	75
Figure 11-5	Standard PM 916, 2006	76
Figure 11-6	Round Robin, Chemex 1 vs. Acme, 2006	78
Figure 11-7	Round Robin, Chemex 1 vs. Chemex 2, 2006	79
Figure 11-8	Round Robin, Acme vs. Chemex 2, 2006	79
Figure 13-1	Results of Grind Tests on New Polaris Bulk Sample	85
Figure 13-2	Locked Cycle Test Schematic Diagram (Process Research Associated Ltd. (PRA), 2007)	98
Figure 13-3	Locked Cycle Test Results (Process Research Associated Ltd. (PRA), 2007)	99
Figure 13-4	Slurry Settling Test Results (Process Research Associated Ltd. (PRA), 2007)	101
Figure 13-5	Slurry Settling Tests on BIOX™ Products (Outotec RS (PTY) LTD, 2018)	104
Figure 13-6	Slurry Settling Test Results on BIOX™ Residues (Outotec RS (PTY) LTD, 2018)	106
Figure 13-7	New Polaris Process Flowsheet	108
Figure 14-1	CPP of Assay Data by Year	112
Figure 14-2	Drillholes used and Domains, 3D orthogonal view looking Northeast	113

Figure 14-3	New Polaris Veins – CPP Plot -Au	114
Figure 14-4	Histogram of Assay Lengths within Mineralized Zone – New Polaris	115
Figure 14-5	Model Extents (black) and the Blocks coded by Majority Domain, 3D Orthogonal View looking Northeast	117
Figure 14-6	Modelled Veins and Underground Workings/Stopes (black), 3D Orthogonal View looking Northeast	120
Figure 14-7	Swath Plots for Au	122
Figure 14-8	Comparison of Assay Au Grades and OK block grades, West East Cross-section, North 870 in mine specific space, looking north	123
Figure 14-9	Comparison of Assay Au Grades and OK block grades, West East Cross-section, North 915 in mine specific space, looking north	124
Figure 14-10	Comparison of Assay Au Grades and OK block grades, North South Cross-section, East 1765 in mine specific space, looking west	125
Figure 16-1	New Polaris Mine Plan Development and Orebody, Sectional View in mine specific space, looking North	130
Figure 16-2	New Polaris Mine Plan Development and Orebody, plan view in mine specific space	131
Figure 16-3	Example of longitudinal Longhole Stoping (from Atlas Copco, 2003)	133
Figure 16-4	Example of Mechanized Cut & Fill (Atlas Copco, 2003)	134
Figure 17-1	New Polaris Process Flowsheet	143
Figure 18-1	CDF in relation to Project	155
Figure 18-2	Conceptual Layout for New Polaris CDF	157
Figure 20-1	Anticipated Regulatory Review Process	162
Figure 22-1	Net Receipts vs. Operating Costs	180
Figure 22-2	Pre-Tax Economic Sensitivity, NPV 5%	180
Figure 22-3	Pre-Tax Economic Sensitivity, IRR	181

The deposit is composed of three sets of veins (quartz-carbonate stringers in altered rock), the “A-B” veins are northwest striking and southwest dipping, the “Y” veins are north striking and dipping steeply east and finally the “C” veins are east-west striking and dipping to the south to southeast at 65º to vertical. The “C” veins appear to hook around to the north and south into the other two sets of veins so that their junctions form an arc. The gold is refractory and occurs dominantly in finely disseminated arsenopyrite grains that mineralize the altered wallrock and stockwork veins. The next most abundant mineral is pyrite, followed by minor stibnite and a trace of sphalerite. The zones of mineralization range from 15 to 250 metres in length and 0.3 to 14 metres in width.

Mr. Robert J. Morris of MMTS conducted a site visit and detailed examination of the property August 22 through to August 23, 2006. During the site visit, sufficient opportunity was available to examine core logging procedures, drill core from the 2006 program as well as conduct a general overview of the property, including selected drill sites, historic core, an underground tour, and the condition of existing project infrastructure. Based on his experience, qualifications and review of the site and resulting data, the author, Mr. Morris, is of the opinion that the programs have been conducted in a professional manner and the quality of data and information produced from the efforts meet or exceed acceptable industry standards. It is also believed that for the most part, the work has been directed or supervised by individuals who would fit the definition of a Qualified Person in their particular areas of responsibility as set out by the Instrument.

The claims are 100% owned and held by New Polaris Gold Mines Ltd., a wholly owned subsidiary of Canarc Resource Corp. (Canarc), and subject to a 15% net profit interest held by Rembrandt Gold Mines Ltd. (Rembrandt), which Canarc has the right to reduce to 10% by issuing 150,000 shares to Rembrandt. Table 4-1 summarizes the claims and the locations are shown on Figure 4-1. Apart from the W.W.1 claim, the claims are crown granted and are kept in good standing through annual tax payments. The W.W.1 is a modified grid claim. The claim has sufficient work filed on it to keep it in good standing until February 4, 2020. The crown granted claims were legally surveyed in 1937. The mineralized areas are shown on Figure 4-2 and Figure 7-2, which shows the geology of the property on the mineral showings.

Claim Name	Lot No.	Folio No.	Claim Name	Lot No.	Folio No.

Polaris No. 1	6109	4472	Snow	3497	4545
Polaris No. 2	6140	5223	Snow No. 2	3495	5088
Polaris No. 3	6141	5223	Snow No. 3	3494	5495
Polaris No. 4	3498	4545	Snow No. 4	3499	5495
Polaris No. 5	6143	5223	Snow No. 5	6105	4472
Polaris No. 6	6144	5223	Snow No. 8	6107	4472
Polaris No. 7	6145	5223	Snow No. 7	3500	4472
Polaris No. 8	6146	5223	Snow No. 6	6106	4472
Polaris No. 9	6147	5223	Snow No. 9	6108	4472
Polaris No. 10	6148	5290	Black Diamond	3491	4472
Polaris No. 11	6149	5290	Black Diamond No. 3	6030	4944
Polaris No. 12 Fr	6150	5290	Blue Bird No. 1	5708	4545
Polaris No. 13 Fr	6151	5290	Blue Bird No. 2	5707	4545
Polaris No. 14	6152	5290	Lloyd	6035	5010
Polaris No. 15	6153	5290	Lloyd No. 2	6036	5010
Silver King No. 1	5489	4804	Rand No. 1	6039	5010
Silver King No. 2	5490	4804	Rand No. 2	6040	5010
Silver King No. 3	5493	4804	Minto No. 2	6033	4944
Silver King No. 4	5494	4804	Minto No. 3	6034	4944
Silver King No. 5	5491	4804	Jumbo No. 5	6031	4944
Silver King No. 6	5492	4804	Ready Bullion	6032	4944
Silver King No. 7	5495	4804	Roy	6042	5088
Silver King No. 8	5717	4545	Frances	6041	5010
Silver Queen No. 1	6026	4545	Eve Fraction	6170	5495
Silver Queen No. 2	6027	4545	Eve No. 1 Fraction	6171	5495
Silver Queen No. 3	6028	4944	P.T. Fraction	3493	5495
Silver Queen No. 4	6029	4944	Ant Fraction	3492	5088
Silver Strand No. 1	6037	5010	Atlin Fraction	3496	5088
Silver Strand No. 2	6038	5010	Powder Fraction	6043	5088
F.M. Fraction	6044	5088	Jay Fraction	6045	5088
Par Fraction	6154	5290

D.C. Sharpstone then secured an option on the property on behalf of Edward C. Congdon and Associates of Duluth, Minnesota. Congdon conducted 236 m (775 feet) of underground exploration in the “AJ” tunnel and collared 26 m (85 feet) into the Canyon adit. The Polaris-Taku Mining Company was then incorporated in 1936 to take over the property from Congdon. Polaris-Taku erected a 150-ton per day flotation mill in 1937 and mined underground continuously until it was closed down in April 1942 due to labor restrictions brought on the Second World War. Mining operations resumed in April 1946 and continued until 1951 when the mine was closed due to high operating costs, a fixed gold price, and the sinking of a concentrate barge shipment during a storm in March 1951. Up to date, 231,604 oz. of gold was produced at a recovered grade of 0.3 oz/ton.

A general distribution of this drilling can be seen in Figure 7-3. Initial efforts were confined to the lower elevations of the property due to limited availability of road building equipment and were designed to test the "Y" Vein system either down dip or along strike from old workings. Discovery of the "C" Vein system in 1989 resulted in a refocusing of efforts towards defining this Zone. Drilling during 1994 and 1995 has been designed to test the North Zone and the downward continuity of the "C" Zone. Drilling on the North Zone cut low-grade gold mineralization in a gently dipping shear zone. Drilling at 60 m (200 foot) centres showed the mineralization to be of limited extent and bounded down dip by a post mineralization fault. No additional drilling of the North Zone is warranted.

Table 6-2 summarizes the variety of estimations identified above by the following: Beacon Hill’s 1988 estimation of residual “reserves” within and around the workings were totaled. To this total, the geostatistical resource estimation of Giroux was added after applying a general dilution factor of 25 % at zero grade to Giroux's figures for the " Y " Zone and 15% at zero grade for the "AB" and "C" Zones. The dilution factors were estimated based on vein characteristics. The "Y" Veins are described as being high grade, but narrow, which makes them prone to high dilution from over-break during mining as well as over mining. The "AB" veins in situ grade, as estimated by Giroux, already contains internal dilution from a parallel dike. To this total, an overall additional dilution of 15 % was added which is appropriate as the "C" vein would not experience much dilution since it is generally thought to be fairly thick. This estimate does not meet the definition requirements of NI 43 – 101 for a resource or reserve. The Author has not done sufficient work to classify them as current reserves or resources and is not treating them as current. This estimate therefore should not be relied upon but is included for historical purposes.

The possible extension of the Llewellyn fault, termed the South Llewellyn fault, continues south from the Chief Cross fault along mine grid coordinate 4400 East. Slightly north of Whitewater Creek it is offset to the west by an east-west fault, the 101 fault, to continue in a more southeast orientation of the opposite side of Whitewater Creek. This northwest-southeast orientation structure was named the Limestone Fault due to its bedding parallel attitude within a discontinuous limestone/marble horizon. It marks the southwest boundary of the “mine wedge”: the wedge-shaped package of rock within which all past production took place. The northern boundary of the “mine wedge” is further defined as mentioned above by the Whitewater Creek Schist Zone, a zone of schistose chlorite-amphibolite-serpentinite less than 100 m thick. A complex network of brittle faults is also found within this zone.

Three major faults, Numbers 1 and 5, and an unnamed fault, lie within the mine wedge. The No.1 and No.5 faults strike northwest-southeast, dipping approximately 45º to the northeast, and are sub-parallel to the unnamed fault, which dips steeply to the southwest. The No.1 fault has reverse displacement of up to 30 m while the displacement of the No.5 fault is poorly defined. The southwest dipping, unnamed fault showed no displacement, as it apparently parallels the A-B vein system. The mined-out areas indicate the wedge shape, the predominant orientations and continuity of the zones, and the overall plunge of the system to the southeast. An early interpretation of the structure showed that various veins appear to meet and form “junction arcs” where both thickness and grade improve.

The vein mineralization has well marked contacts with the wall rock. The transition from mineralized to non-mineralized rock occurs over a few centimeters. The mineralization consists of at least three stages of quartz veining. The initial stage of quartz-ankerite introduced into the structure was accompanied by a pervasive hydrothermal alteration of the immediately surrounding wall rock. Arsenopyrite, pyrite and lesser stibnite were deposited with the alteration. Later stages of quartz-ankerite veining are barren and have the effect of diluting the gold grades in the structure. The sulphide minerals are very fine-grained and disseminated in both the wall rock and early quartz and ankerite veins. Free gold is extremely rare and to the end of 2005 had not been recognized in core samples. The majority of the gold occurs in arsenopyrite and to a lesser extent in pyrite and stibnite. Because there is no visible gold and the host sulphides are very fine-grained and disseminated there is little nugget effect and gold values even over short intervals rarely exceed 1 oz/ton.

In 1996 and 1997 the Y, C and AB veins were explored from underground. The plan was to closely test the upper portions of the Y, C and AB veins in order to allow calculation of a resource that might form the basis for resumption of mining. The results of the underground drilling program were mixed. The underground workings were for the most part driven along the vein structures with few crosscuts from which holes could be drilled to cut the down dip and along strike extension of the veins. As a result, except for those holes that tested the area immediately below the workings, most cut the veins at shallow angles. The very shallow angles that in places approach parallel to the vein make the use of these intersections inappropriate for a resource calculation (An example is hole 97-44 that cut 34.1 m grading 0.42 oz/ton). Despite the number of holes drilled during 1996 and 1997, the work did little to expand the extent of the mineralization in the AB, C or Y vein systems. The work did confirm that the mineralized shoots in the lower most stopes on the Y and C veins were open to depth. The Table below summarizes the locations of the 1988 to 2005 drilling. Composites of assay results are listed for the C vein system holes that intersected significant mineralization.

The results of the 2003 to 2006 drilling of the C vein system confirmed the continuity of gold mineralization and the vein structure between the earlier drilled holes. As can be seen in the sections below, drill results show the C vein system to be an arc-like structure oriented east-west in the west swinging to a northeastern strike in the east. The change in strike occurs across the No.1 fault. To the east of the No.1 fault, the vein splays into two or more branches. The dip of the vein system is to the south and southeast and has an average dip of about 50º, although east of the No.1 fault the vein appears to flatten and thicken in a simoid-like feature. The exact nature of the apparent flattening of the vein’s dip is not clear and requires additional drilling to be resolved.

04-1737E2	1737.33	845.82	289.26	-83	320
04-300SW1	1842.61	902.79	224.03	-66.1	320
04-300SW2	1841.57	901.6	246.89	-75	320
04-300SW3	1842.7	902.73	248.41	-83	320
04-330SW1	1785.71	922.54	213.06	-81.7	323.8
04-330SW2	1789.62	904.06	237.44	-83.8	324
04-360SW1	1777.39	884.53	237.44	-75	318.5
04-360SW2	1777.39	884.53	243.84	-83	318
05-1676E1	1676.4	872.49	210.92	-71	359
05-1676E2	1676.4	872.49	240.49	-78	359
05-1676E3	1676.4	872.49	94.49	-83	359
05-300SW4	1875.64	860.29	289.86	-77.5	322
05-300SW5	1877.2	858.36	315.16	-85	318
05-300SW6	1911.06	816.86	347.17	-84.5	320
05-330SW3	1835.64	860.95	277.98	-77	320
05-330SW4	1835.69	860.88	298.4	-83	320
05-330SW5	1853.22	839.92	315.16	-83	320
L52	1465.17	1330.15	248.41	-25	85
L55	1510.89	1361.54	274.32	-17	40
L56	1498.7	1313.69	161.85	-25	92
L57	1464.26	1330.45	215.49	-25	51
L58	1473.4	1278.79	189.28	-25	112
L59	1477.67	1279.25	343.2	-25	95
L62	1463.04	1233.83	306.32	-25	149
L84	1342.64	1237.79	229.21	-2.5	247.5
P-9612	1289.97	1259.92	174.65	-35	119
P8918A	1819.79	889.97	367.89	-70	315
P91C01	1888.57	762.3	322.17	-65	338
P91C02	1888.57	762.3	331.62	-77	338
P91C03	1937.98	805.62	320.04	-75	339
P91C04	1937.98	805.62	324.92	-87	339
P91C05	1827.28	942.78	199.19	-73	339
P91C06	1672.83	818.27	263.96	-71	339
P91C07	1907.53	704.18	320.65	-75	335
P91Y01	1767.03	1369.25	330.1	-66	270
P91Y02	1769.59	1336.46	305.1	-64	266
P91Y03	1705.02	1095.33	274.93	-49	268
P91Y04B	1845.38	1336.03	444.09	-68	274
P92C08	1810.12	984.99	199.03	-71	342
P92C09	1889.24	964.2	210.31	-72	340
P92C10	1917.31	894.37	243.84	-76.5	340
P92C11	1978.82	925.86	251.16	-71.5	337
P92C12	1870.25	819	282.55	-71.5	337
P92C13	1635.16	860.39	223.72	-70	345
P92C14	1614.98	823.9	262.13	-70	343
P92C15	1702.92	823.57	258.17	-71.5	338
P92C16	1816	780.29	254.2	-71.5	338
P92C17	1730.41	865.11	264.26	-76	339
P92C18	1606.63	792.6	350.22	-70	337
P92C19	1694.87	761.73	381.3	-70	350
P92C20	1549.57	832.74	267.31	-70	350
P92C20A	1549.57	832.74	273.71	-70	350
P92C21	1534.49	871.36	213.97	-71	358
P92C22	1624.58	743.71	413.61	-71	336
P92C23	1518.98	873.06	274.62	-72	337
P92C24	1798.02	918.21	149.66	-61	331
P92Y05	1769.67	1335.82	346.56	-70	259
P92Y08	1787.35	1367.03	326.96	-67	274
P92Y09	1745.28	1336.55	248.11	-61	269
P93C25	1737.36	911.35	102.41	-65.5	340
P93C26	1858.98	1036.32	144.17	-70	295
P93C27	1860.13	1034.89	166.73	-85	295
P93C28	1882.41	1050.04	165.51	-75	310
P93C29	1906.65	1088.81	179.83	-75	280
P93C30	1928.77	1144.68	168.25	-75	280
P93C31	1815.61	901.17	215.49	-74	335
P94C32	1908.04	962.37	249.02	-72	303
P94C33	1908.37	961.28	279.81	-86	284
P94C34	1937.31	1015.9	258.17	-72	287
P94C35	1937.31	1015.9	339.55	-83	288
P94C36	1957.12	891.54	371.86	-80	288
P94C37	1883.66	789.74	279.5	-74	327

P94C38	1699.87	806.2	287.73	-82	348
P94C39	1769.36	768.1	290.17	-77	351
P94N01	1469.75	2124.76	84.73	-55	85
P94N02	1469.75	2124.76	84.73	-85	85
P94N03	1488.95	2194.26	60.35	-55	85
P94N04	1488.95	2194.26	106.07	-90	0
P94N05	1495.96	2258.26	118.26	-56	85
P94N06	1495.96	2258.26	124.36	-87	85
P94N07	1488.34	2332.63	224.94	-55	85
P94N08	1488.34	2332.63	273.71	-85	85
P94N09	1490.17	2438.1	264.57	-50	70
P94N10	1481.33	2438.1	96.93	-85	70
P94N11	1492.3	2499.36	75.59	-50	70
P94N12	1492.3	2499.36	71.02	-85	70
P94N13	1482.85	2384.76	66.45	-50	70
P94N14	1482.85	2384.76	66.45	-85	70
P94N15	1187.2	1822.7	51.21	-60	70
P94N16	1187.2	1822.7	57.3	-60	100
P94Y10	1663.9	1203.96	236.83	-63	266
P94Y11	1813.56	1280.16	251.76	-55	270
P94Y12	1813.56	1280.16	276.15	-63	265
P94Y13	1929.69	1191.46	371.86	-62	265
P94Y14	1929.69	1191.46	400.81	-55	262
P95C40	1965.96	575.77	786.69	-82.5	355
P95C41	2029.97	553.21	769.01	-80	360
P95C42	1861.72	571.5	673	-77.5	0
P95C43	1861.72	571.5	751.94	-80	340
P95C44	1920.24	525.78	792.48	-83	340
P95N17	1347.22	2122.93	264.57	-69.5	85
P95N18	1370.69	2183.89	270.66	-72	85
P95N19	1350.26	2243.33	315.47	-68	85
P95N20	1348.74	2302.76	335.28	-70	85
P95N21	1348.13	2329.59	215.8	-70	70
P95N22	1353.92	2387.8	227.99	-72	70
P95N23	1338.07	2449.07	261.52	-68	70
P95N24	1336.24	2449.07	294.74	-85	70

P95N25	1502.66	2804.16	255.42	-55	100
P95N26	1325.88	2804.16	273.71	-80	100
P95N27	1325.88	2804.16	334.67	-80	80
P9705A	1055.52	1545.03	231.34	-21	52
PC25A	1738.27	909.83	159.41	-68	336
PC44A	1920.24	525.78	749.81	-83	340
PC8801	1554.16	1178.32	106.68	-50	310
PC8802	1554.16	1178.32	38.71	-90	0
PC8803	1496.69	1146.74	133.2	-60	310
PC8804	1496.21	1141.91	119.48	-60	280
PC8805	1614.15	1363.59	149.35	-45	271
PC8806	1614.68	1363.74	189.28	-60	270
PC8807	1614.14	1364.2	180.14	-48	290
PC8808	1614.26	1365.03	110.95	-48	310
PC8901	1573.08	1149.96	142.34	-90	0
PC8902	1594.63	1088.77	128.32	-86	275
PC8903	1593.56	1088.51	167.64	-55	265
PC8904	1546.21	1068.5	50.9	-52	90
PC8904A	1546.21	1068.5	230.73	-52	90
PC8905	1546.73	1039.04	167.94	-60	90
PC8906	1705.42	1363.04	220.98	-55	268
PC8907	1704.52	1363.2	236.83	-70	290
PC8908	1704.52	1363.2	243.23	-50	290
PC8909	1639.71	1341.13	185.32	-45	270
PC8910	1639.14	1341.19	126.34	-59	270
PC8911	1639.77	1342.07	187.45	-53	310
PC8912	1726.56	1440.13	307.85	-45	270
PC8913	1727.76	1440.48	244.75	-60	270
PC8914	1726.95	1439.69	253.59	-56	256
PC8915	1726.18	1163.72	413.31	-65	270
PC8916	1762.08	1377.37	233.78	-55	270
PC8917	1676.68	1157.69	171.3	-45	270
PC8918	1819.79	889.97	191.11	-70	315
PC8919	1777.26	1103.56	219.15	-60	270
PC9001	1820.17	822.27	400.51	-70	340
PC9002	1838.2	767.83	388.92	-70	340

PC9003	1746.57	837.7	305.1	-70	340
PC9004	1853.47	897.63	305.71	-70	340
PC9004A	1853.47	897.63	108.91	-70	340
PC9005	1675.88	852.91	265.48	-60	340
PC9006	1762.29	792.22	404.16	-70	340
PC9007	1605.02	880.79	242.62	-68	340
PC9008	1778.45	746.71	441.05	-70	330
PT23A	1317.96	1244.5	206.65	-32	195
PT29A	1317.96	1244.5	325.53	-31	134
PT30A	1317.96	1244.5	279.81	-15	90
PT31A	1334.41	1456.94	406.3	0	20
PT37B	1595.05	1102	179.53	-52	45
PT38A	1595.32	1097.49	260.3	-35	124
PT9601	1444.26	1379.06	106.07	-36	79
PT9602	1441.21	1379.06	120.4	-51	79
PT9603	1441.22	1378.72	102.11	-36	96
PT9604	1441.22	1378.72	130.76	-51	96
PT9605	1441.09	1379.36	101.8	-49	68
PT9606	1441.33	1378.29	115.82	-38	110
PT9607	1436.92	1380.33	43.59	21.5	291
PT9608	1436.92	1380.62	17.98	21.5	299
PT9609	1437.02	1380.72	119.48	21.5	304
PT9610	1437.02	1380.73	119.48	-4.5	304
PT9611	1436.91	1380.62	125.88	-22	299
PT9613	1289.76	1259.71	149.05	-41	134
PT9614	1289.43	1258.88	121.01	-52	164
PT9615	1289.76	1259.71	147.52	-52	134
PT9616	1289.97	1259.92	176.78	-46	119
PT9617	1288.85	1263.64	152.1	0	15
PT9618	1055.52	1545.03	146	18	88
PT9619	1055.52	1545.03	127.41	18	108
PT9620	1055.52	1545.03	173.74	-10	118
PT9621	1055.52	1545.03	141.43	-8.5	80
PT9622	1055.52	1545.03	152.1	22	108
PT9623	1055.52	1545.03	167.34	-14	110
PT9624	1055.52	1545.03	272.19	-10	52

PT9701	1055.52	1545.03	9.75	-5	52
PT9702	1055.52	1545.03	287.73	-7	52
PT9703	1055.52	1545.03	220.07	-13	60
PT9704	1055.52	1545.03	205.44	-10	60
PT9706	1055.52	1545.03	255.42	-9.5	49
PT9707	1055.52	1545.03	240.18	10	52
PT9708	1049.12	1542.59	144.48	0	242
PT9709	1217.37	1301.04	69.8	-22	0
PT9710	1217.98	1301.19	69.49	-22	340
PT9711	1218.9	1301.04	163.37	-15.5	78
PT9712	1218.9	1301.04	171.3	4	60
PT9713	1222.1	1299.97	281.33	-15	90
PT9714	1219.2	1297.84	47.61	-42	0
PT9715	1218.9	1301.04	59.74	-46	330
PT9716	1217.07	1296.62	135.33	-15	270
PT9717	1217.07	1296.62	154.84	-30	270
PT9718	1217.07	1296.62	118.26	-18	249
PT9719	1290.83	1417.62	112.47	-5	90
PT9720	1290.83	1417.62	133.81	-35	90
PT9721	1290.52	1417.62	153.92	-44	71
PT9722	1289.3	1418.23	219.15	-5	270
PT9723	1317.96	1244.5	17.98	-32	195
PT9724	1317.96	1244.5	192.94	-30	217
PT9725	1317.96	1244.5	244.14	-47	218
PT9726	1317.96	1244.5	124.66	-10	334
PT9727	1317.96	1244.5	110.34	-12	351
PT9728	1317.96	1244.5	228.6	-31	351
PT9731	1334.11	1456.94	19.2	0	20
PT9732	1594.99	1102.89	229.51	-22	26.5
PT9733	1594.65	1102.28	272.49	-52	23.5
PT9734	1595.23	1101.76	155.14	-32	50
PT9735	1594.96	1102.83	198.73	-10	26
PT9736	1594.87	1102.95	237.13	-29	24
PT9738	1595.35	1097.49	14.33	-35	124
PT9739	1594.96	1097.43	204.52	-29	136
PT9740	1594.53	1097.25	180.75	-34	146

Hole-ID	From	To	Length (m)	Au_Comp (g/t)
04-1707E1	201.93	203.03	1.10	10.20
04-1707E1	209.46	213.42	3.96	5.10
04-1707E2	222.08	224.73	2.65	10.10
04-1707E2	246.64	248.72	2.07	11.80
04-1737E1	184.25	190.80	7.77	17.13
04-1737E1	200.25	207.57	7.32	4.00
04-1737E1	231.19	235.37	4.18	13.60
04-1737E2	233.66	233.96	0.30	30.90
04-1737E2	256.64	259.08	2.44	9.20
04-300SW1	180.75	181.72	0.98	11.20
04-300SW2	181.20	181.66	0.46	5.50
04-300SW2	206.53	208.94	2.41	14.80
04-300SW3	196.38	203.61	7.22	7.20
04-300SW3	223.88	230.73	6.86	14.50
04-330SW1	167.18	168.80	1.62	4.00
04-330SW1	172.36	174.49	2.13	35.30
04-330SW2	168.86	171.60	2.74	5.40
04-330SW2	194.22	202.27	8.05	31.90
04-360SW1	162.76	163.67	0.91	5.10
04-360SW1	179.10	193.43	14.33	11.60
04-360SW2	191.20	195.38	4.18	25.70
04-360SW2	198.24	201.87	3.63	5.20
05-1676E1	185.62	191.20	5.58	12.10
05-1676E2	210.01	213.27	3.26	8.90
05-300SW4	200.13	200.86	0.73	13.40
05-300SW4	261.34	269.05	7.71	17.40
05-300SW5	215.80	218.30	2.50	14.10
05-300SW5	241.10	246.74	5.64	21.60
05-300SW6	233.54	240.79	7.25	16.70
05-300SW6	260.91	266.09	5.18	18.20

05-330SW3	213.94	223.11	9.17	9.90
05-330SW4	221.89	233.00	11.89	19.88
05-330SW5	230.89	246.28	15.39	8.57
06-1676E-6	284.8	292.2	7.4	8.0
06-1707E-3	232.7	235.8	3.1	17.4
06-1707E-3	247.0	250.0	3.0	5.9
06-1707E-3	253.0	254.9	1.9	10.5
06-1707E-6A	312.0	313.4	1.4	18.8
06-1737E-3	242.3	244.3	2.0	24.6
06-1737E-3	252.7	254.7	2.0	10.6
06-240SW-8	263.5	270.0	6.5	7.7
06-240SW-8	297.7	299.2	1.5	8.8
06-240SW-8	316.1	317.55	1.45	8.3
06-270SW-4	263.1	268.85	5.75	12.3
06-270SW-4	310.8	313.7	2.9	4.9
06-1615E-8	346.4	352.6	6.2	44.7
06-1768E-2	259.2	262.7	3.5	17.9
06-1768E-3	329.6	332.1	2.5	9.1
06-270SW-2	201.6	207.5	5.9	23.0
06-270SW-3	239.5	252.15	12.65	7.4
06-270SW-3	262.0	264.0	2.0	17.5
06-330SW-11	361.0	363.5	2.5	6.6
06-1768E-1A	225.8	242.3	16.5	23.1
06-1646E-6	258.3	265.5	7.2	15.5
06-1676E-5B	257.2	271.2	14.0	10.9
06-300SW-8	303.3	338.3	35.0	8.9
06-330SW-9	326.4	333.0	6.6	8.3
06-330SW-10	330.5	333.25	2.75	12.9
06-330SW-10	373.85	376.45	2.6	17.7
06-1585E-8	391.2	392.05	0.85	7.7
06-1615E-2	142.1	147.3	5.2	25.0
06-1615E-7.5	351.15	354.2	3.05	22.5
06-1615E-9	439.5	451.9	12.4	16.1
06-1813E-1	300.1	303.1	3.0	16.7
06-1813E-2	313.3	314.9	1.6	12.7
06-1813E-2	331.5	333.0	1.5	6.4
06-1813E-2	336.3	343.5	7.2	18.2
06-1813E-2	338.2	339.1	0.9	38.4
06-1813E-3	332.6	333.6	1.0	6.3
06-1813E-3	392.5	394.35	1.85	6.0
06-1813E-3	405.7	407.25	1.55	7.3
06-1707DE-1	425.6	430.1	4.5	11.8
06-1768DE-1	473.3	474.0	0.7	12.6

06-1768DE-1	487.4	491.9	4.5	9.5
06-1768DE-2	538.4	548.6	10.2	7.1
06-1859E-2	297.5	299.1	1.6	14.9
06-1859E-2	336.2	338.3	2.1	15.6
06-1859E-2	417.4	420.8	3.4	11.0
P8918A	194.43	199.49	5.06	37.70
P91C01	251.83	258.17	6.34	7.80
P91C01	274.93	281.48	6.55	20.20
P91C02	263.35	268.10	4.75	25.70
P91C03	226.47	229.64	4.20	11.46
P91C04	282.55	288.74	6.19	4.80
P91C05	156.73	169.10	12.37	10.30
P91C06	228.66	231.31	2.65	22.40
P91C07	292.06	294.35	2.29	18.80
P92C08	124.05	127.35	3.29	8.90
P92C09	161.85	163.98	2.13	4.20
P92C12	216.26	218.08	1.83	4.20
P92C12	258.47	265.27	6.80	15.20
P92C13	190.56	192.39	1.83	15.00
P92C14	246.10	248.69	2.59	6.90
P92C15	231.77	233.17	1.40	4.80
P92C15	236.89	238.66	1.77	18.70
P92C16	229.36	231.65	2.29	4.20
P92C17	209.70	213.48	3.78	21.80
P92C18	281.18	283.98	2.80	9.20
P92C19	282.64	285.51	2.87	23.80
P92C20	243.35	245.97	2.62	8.90
P92C20A	233.08	235.82	2.74	12.50
P92C21	172.15	178.06	5.91	18.70
P92C24	136.52	139.39	2.87	23.70
P93C26	108.02	110.95	2.93	26.60
P93C27	148.89	151.49	2.59	12.50
P93C28	140.27	142.43	2.16	10.50
P93C31	184.80	186.99	2.19	8.00
P94C33	266.43	269.53	3.11	6.90
P94C34	176.72	180.14	3.41	7.20
P94C36	261.21	262.28	1.07	6.40
P94C37	255.24	261.27	6.04	22.20
P94C38	255.54	257.13	1.58	20.10
P94C38	274.53	279.23	4.69	22.90
P94C39	262.01	266.58	4.57	28.70
P95C40	493.01	498.93	5.92	8.47
P95C40	727.77	735.24	7.47	11.31

Three tests were completed to compare the duplicate sample results: a Student’s t-test, which is a comparison of the mean values; an F-test, which is a comparison of the variance; and a Paired t-test, which is a test for bias. The Student’s t-test shows that the means of the duplicate samples are likely to be the same (the calculated t-value is -0.32, well below the critical value of 1.96). The F-test shows that the variances of the two sample sets are likely to represent the same population (the calculated F-value is 0.61, within the range 0.065 and 1.68 of F-values characteristic of a normal distribution). The Paired t-test shows that the difference between the two data sets is minimal, and they are distributed around the value zero (the standard error is calculated to be 0.18, while the mean, 0.30, is between –0.065 and 0.657).

Three tests were completed to compare the round robin sample results: a Student’s t-test, which is a comparison of the mean values; an F-test, which is a comparison of the variance; and a Paired t-test, which is a test for bias. The Student’s t-test shows that the means of the round robin samples are likely to be the same (the calculated t-value is -0.18 for the Chemex 1 vs. Acme samples; 0.11 for the Chemex 1 vs. Chemex 2 samples; and 0.07 for the Acme vs. Chemex 2 samples, well below the critical value of 1.96). The F-test shows that the variances of the three sample sets are likely to represent the same population (the calculated F-value is 1.04, for the Chemex 1 vs. Acme samples; 1.12 for the Chemex 1 vs. Chemex 2 samples; and 0.96 for the Acme vs. Chemex 2 samples, within the range 0.065 and 1.68 of F-values characteristic of a normal distribution). The Paired t-test shows that the difference between the Acme and Chemex 2 data set is minimal, and they are distributed around the value zero (the standard error is calculated to be 0.08, while the mean, 0.19, is between –0.065 and 0.657). In both cases, the Chemex 1 vs. Acme and Chemex 1 vs. Chemex 2 data sets, show a slight bias with the Chemex 1 samples being 0.32 to 0.33 g/t lower than the Chemex 2 and Acme respectively.

Laboratory	Sample Type	Tests	Comments
Resource Development Inc. (RDI) (2003)	Bulk sample	Grinding	Grind tests indicated a P80 of 75 lm.
		Diagnostic Leach	Indicated the following gold distribution: 66.4% Arsenopyrite/Pyrite, 20.1% Quartz, 9% Free gold, 4.5% Stibnite gold.
		Rougher Flotation	Overall good results observed (83 – 94% recovery).
		Rougher Concentrate Production – Test 4	17.6% mass pull, 90.15% AU recovery, approximately 15% carbonates.
	Rougher Concentrate from Test 4	Cleaner Flotation	Indicated only 1 stage of cleaning required. Cleaner flotation not sensitive to residence time. Test 7: Rougher test to generate concentrate for Tests 8, 9.18.9% mass pull with 96.1% Au recovery. Confirmed no regrind required prior to cleaning.
	Tails and Conc. From Test 4	Gravity Concentration	Indicated ore is generally not amenable to gravity concentration.
	Rougher and cleaner Tails	Cyanidation – Bottle Roll	Poor gold recovery, high NaCN consumption.
	Bulk Sample	Rougher Concentrate Production	19.5 initial wt., 96.6% Au recovery, 12 kg conc. Concentrate shipped to Mintek and Oxidor, tested by RDI.
Mintek(2003)	Concentrate from RDI Rougher Test (2003)	Cyanidation – CIL and Nitric Acid pre-treatment.	17.6% extraction vs. 93.7% after nitric acid treatment. Confirms sulphide oxidation required for gold recovery.
		Diagnostic Leach	Confirmed majority gold content is refractory in sulphides.
		Bio Leach	2 Tests: normal bio-leach, bio-leach + acetone and ferric pre-wash. Poor dissolution, indicated inhibitory substance (As+3 or As+5).
RDI (2004)		Batch Acid Pressure Oxidation (POX)	100% Sulphur oxidation after 1.5, 1 hour, respectively.
RDI (2004)		Cyanidation – Bottle Roll Tests	9.4% baseline Au recovery, > 98% Au recovery after POX (1, 1.5 hours).
Oxidor Laboratory(2004)		Bio-Oxidation (BIOX™) Amenability Tests	Bio-oxidation test using OXL-1014-R-13 culture; adaptation, inoculation, build-up.98% sulphide oxidation, 90% Au extraction after 9 days BIOX™. Confirmed inhibitory substance (As+3 or As+5).
Oxidor Laboratory(2004)		Cyanidation – CIL	8.2% Au recovery without BIOX™, compared to 90% Au extraction after BIOX™.
Process Research Associates Ltd. (PRA)(2007)	Bulk Sample	Cyanidation – Bottle Roll Tests	Low Au recovery.
		Gravity Separation	Confirmed not amenable to gravity.
		Rougher and Cleaner Flotation	Tested various flotation conditions for flotation optimization. Pyrite/Arsenopyrite separation unsuccessful. Best results achieved after 1 rougher, 1 scavenger, and 1 cleaner stage: 15.2% mass pull, 94.9% Au recovery.
		Locked Cycle Flotation	Multiple recycled streams used, 5 cycles with 3 stages of cleaning.
	Final bulk flotation tails	Slurry Settling Test	Two settling tests: with/without Percol 156 flocculant. 21% solids increased to 71% solids, 0.9 m/h settling rate increased to 2.5 m/h with flocculant.
	Flotation Concentrate	Cyanidation – High Intensity Leach	Poor results: 10.6% Au recovery, 31 kg/t NaCN consumption.
Outotec RSA (PTY) Ltd. (2018)	Flotation Conc.	BIOX™ Batch Amenability Tests (BAT)	Inoculum Adaptation, 7 BATs (12 – 22-day oxidation times). Achieved 89.6 – 99.1% Sulphide Oxidation.
	Flotation Conc. and BIOX™ Residue	Cyanidation – Bottle Roll Tests	8.1% baseline Au recovery. Approximately 95.7% Au recovery after 22-day BIOX™.
	BIOX™ Liquor from BAT 1	BIOX™ Liquor Neutralization Test (2018)	3 Tests: lime/limestone, lime only, and lime/limestone + Fe2(SO4)3. Slow neutralization, [Fe], [As] indicates As can be reduced to below EPA limit of 0.4 mg/L. Long term As stability requires addition of Fe2(SO4)3.
	BIOX™ Slurry from BAT 3	BIOX™ Residue Static Settling Tests (2018)	8 flocculants tested for clarity of liquor, dosage tests once flocculant chosen. 150 – 250 g/t Magna 405 flocculant produces good settling rate, 300 g/t, may improve results slightly.

Test #	Comments	Float time (min)	Recovery (% of feed)				Assay Value
Test #	Comments	Float time (min)	Wt.	Au	As	CO2	Au (g/t)	As (%)	CO2 (%)
5	Cleaner 2 Con 1	5	43.4	63.2	62.1	3.1	118.00	15.00	0.54
	Cleaner 2 Con 2	4	11.8	22.7	23.5	7.3	155.14	20.80
	Cleaner 2 Tail		5.9	7.1	7.4	5.2	98.68	13.10	6.58
	Calc. Cleaner 1 Conc.	10	61.1	93.0	93.0	15.6	123.3	15.94
	Cleaner 1 Tail		38.9	7.0	7.0	84.4	14.54	1.89	16.20
6	Cleaner 1 Con 1	1	42.1	62.0	63.2	6.5	122.02	16.10	1.10
	Cleaner 1 Con 2+3	9	20.6	32.5	31.5	14.9	130.80	16.40	5.20
	Cleaner 1 Tail		37.3	5.5	5.3	78.6	12.07	1.52	15.10
7	Rougher Conc.	24	18.9	96.1	97.4		93.27	13.08
7	Rougher Tail		81.1	3.9	2.6		0.89	0.08
8	Cleaner Conc.	10	80.8	97.2	97.9		106.78	15.7
8	Cleaner Tail		19.2	2.8	2.1		13.17	1.44
9	Cleaner Conc.	8	46.7	56.5	55.2		116.26	15.60
9	Cleaner Tail		53.3	43.5	44.8		78.46	11.10

Test #	Test Conditions	Conc. or Tails	Recovery (%)				Grade (g/t, %)
Test #	Test Conditions	Conc. or Tails	Au	As	Sb	ST	Au (g/t)	As (%)	Sb (%)	ST (%)
F6	pH 8.9Rougher: Na2SiO3, 10 g/t A3418, 50 g/t PAX, 19 g/t MIBCScavenger (x3): 15 g/t A3418, 75 g/t PAX, 9 g/t MIBC20 minute regrindCleaner (x3): 650 g/t Na2SiO3, 10 g/t A3418, 51 g/t PAX, 9 g/t MIBC	Cleaner Conc.	45.3	37.7	51.2	66.7	63.2	8.27	5.99	29.7
		Cleaner Tails	24.7	23.9	6.6	11.4	89.0	13.5	1.99	13.1
		Ro/Scav Bulk	91.6	91.6	71.0	93.2	25.1	3.95	1.63	8.14
F7	pH 8.8Rougher: Na3PO4, 10 g/t A3418, 50 g/t PAX, 19 g/t MIBCScavenger (x3): 15 g/t A3418, 75 g/t PAX, 9 g/t MIBC20 minute regrindCleaner x3): 650 g/t Na3PO4, 10 g/t A3418, 51 g/t PAX, 9 g/t MIBC	Cleaner Conc.	47.1	37.6	55.1	65.1	65.2	7.08	6.52	28.9
		Cleaner Tails	16.9	18.8	4.5	7.2	100.2	15.2	2.29	13.8
		Ro/Scav Bulk	90.0	88.5	73.3	93.2	25.4	3.40	1.77	8.45
F8	Grind: 250 g/t Na2SRougher: 50 g/t PAX, 13 g/t MIBCScavenger (x2): 50 g/t PAX, 13 g/t MIBC, 250 g/t CuSO4Cleaner (x2): 50 g/t Na2SiO3, 10 g/t PAX, 15 g/t MIBC	Cleaner Conc.	94.9	89.9	91.8	94.5	85.68	8.98	4.65	18.97
		Cleaner Tails	1.8	4.4	4.0	2.1	2.20	0.59	0.27	0.57
		Ro/Scav Bulk	96.7	94.3	95.8	96.6	49.96	5.39	2.78	11.10

Test #	Test Conditions	Conc. or Tails	Mass Pull (%)	Recovery (%)				Grade (g/t, %)
Test #	Test Conditions	Conc. or Tails	Mass Pull (%)	Au	As	Sb	ST	Au (g/t)	As (%)	Sb (%)	ST (%)
F8	Grind: 250 g/t Na2S Rougher: 50 g/t PAX, 13 g/t MIBC Scavenger (x2): 50 g/t PAX, 13 g/t MIBC, 250 g/t CuSO4 Cleaner (x2): 50 g/t Na2SiO3, 10 g/t PAX, 15 g/t MIBC	2nd Cleaner Conc.	12.8	93.8	87.4	88.9	93.3	100.61	10.37	5.35	22.26
		2nd Cleaner Tails	2.4	1.1	2.5	2.8	1.2	6.32	1.56	0.90	1.53
		1st Cleaner Conc.	15.2	94.9	89.9	91.8	94.5	85.68	8.98	4.65	18.97
		1st Cleaner Tails	11.3	1.8	4.4	4.0	2.1	2.20	0.59	0.27	0.57
		Rougher Bulk	26.5	96.7	94.3	95.8	96.6	49.96	5.39	2.78	11.10
		Rougher Tails	73.5	3.3	5.7	4.2	3.4	0.61	0.12	0.04	0.14
		Total	100	100	100	100	100	9.70	2.02	0.78	3.18

Outotec conducted a testing campaign to determine the amenability of bio-leaching the New Polaris flotation concentrate using the Outotec BIOX™ technology. Tests were conducted using 250 g of material (80% passing 75 lm) in a “9K” nutrient broth at 7.7% solids at pH 1.5 (with sulphuric acid). The samples were then inoculated with a bacterial culture (applied at 10% solids) and kept between 38 – 42°C at a pH of 1.2 – 1.4. Bacterial activity was then monitored, and more ore was added once activity increased to desirable levels until 20% solids was achieved in the conditioning reactor. Once the ferrous iron levels reduced to 0.1 g/L in solution, 300 mL of the slurry was extracted and transferred to a 3 L beaker with “0K” nutrient broth, and 20% solids was then achieved for the build-up phase.

Assay sample lengths have varied with the drill programs. A histogram of the assay intervals for the New Polaris deposit is shown in Figure 14-4. The majority of sampling has been done using a 1.5 m interval length. Therefore, this is the value that has been used for the base length when compositing. The compositing also honored the domain boundaries. Assay intervals less than 0.75 m have been added to the previous composite. To ensure correct compositing the length weighted mean grades of the composite are compared to the original assay data. Assay and composite statistics are summarized in Table 14-5 showing negligible differences in grade values between composites and assays coded with a domain. Theses statistic also indicated that the Coefficient of Variation (C.V.), at values primarily around 1.0 and not over 1.85 are appropriate for using a linear interpolation method.

Source	Parameter	Vein
Source	Parameter	1	2	5	6	7	8	ALL
Assays	Num Samples	186	440	450	191	133	43	1443
	Num Missing Samples	7	7	1	5	1	0	21
	Min (g/t)	0	0	0	0	0	0	0
	Max (g/t)	33.02	80.61	118.50	69.60	111.94	45.94	118.50
	Weighted mean (g/t)	6.71	10.73	11.28	4.96	6.72	12.03	9.18
	Weighted SD	6.35	13.04	14.07	10.18	14.61	12.68	12.79
	Weighted CV	0.95	1.22	1.25	2.05	2.18	1.05	1.39
Comps	Num Samples	110	282	292	139	100	32	955
	Num Missing Samples	21	10	3	6	0	0	40
	Min	0	0	0	0	0	0	0
	Max	27.60	56.52	118.50	57.07	57.70	42.23	118.50
	Weighted mean	6.72	10.66	11.28	4.91	6.68	12.03	9.14
	Weighted SD	5.43	10.89	11.78	9.11	10.95	11.11	10.76
	Weighted CV	0.81	1.02	1.04	1.85	1.64	0.92	1.18
Mean Au Grade Difference (%)		0.2%	-0.6%	0.0%	-1.0%	-0.5%	0.0%	-0.4%

Domain	Azimuth	Dip	Axis	Total Range (M)	NUGGET	SILL 1	SILL 2	Short Range	Long Range
Domain	Azimuth	Dip	Axis	Total Range (M)	NUGGET	SILL 1	SILL 2	(m)	(m)
11	ROT	70	Major	100	0.3	0.4	0.3	20	100
	DIPN	0	Minor	120				40	120
	DIPE	-45	Vertical	90				10	90
21	ROT	70	Major	70	0.2	0.7	0.1	40	70
	DIPN	0	Minor	90				50	90
	DIPE	-45	Vertical	60				40	60
51 - EAST	ROT	110	Major	60	0.4	0.3	0.3	20	60
	DIPN	-10	Minor	70				15	70
	DIPE	-70	Vertical	40				10	40
52 - WEST	ROT	110	Major	60	0.4	0.3	0.3	20	60
	DIPN	-10	Minor	70				15	70
	DIPE	-80	Vertical	40				10	40
61	OMNI DIRECTIONAL			60	0.6	0.3	0.2	10	60
71	ROT	0	Major	80	0.4	0.4	0.2	60	80
	DIPN	0	Minor	70				30	70
	DIPE	-85	Vertical	30				10	30
81	ROT	340	Major	80	0.4	0.4	0.2	60	80
	DIPN	0	Minor	70				30	70
	DIPE	-90	Vertical	30				10	30
Waste	OMNI DIRECTIONAL			80	0.7	0.2	0.1	20	80

Source	Au Grade	Estimation Domain
Source	Parameter	11	21	51	52	61	71	81	ALL
Model Values – Ordinary Kriging	Num Samples	3243	5649	1545	4105	946	844	291	16623
	Num Missing Samples	3	0	0	3	0	0	0	6
	Min	0.049	0.019	0.163	0.951	0.000	0.005	0.062	0.000
	Max	20.94	38.94	43.73	35.97	26.72	30.69	11.93	43.73
	Weighted mean	8.68	14.38	16.97	13.01	8.99	13.89	6.92	13.31
	Weighted CV	0.42	0.46	0.65	0.38	0.54	0.63	0.43	0.52
De-Clustered Composites	Num Samples	3243	5649	1545	4105	946	844	291	16623
	Num Missing Samples	3	0	0	3	0	0	0	6
	Min	0.000	0.000	0.000	0.000	0.000	0.000	0.000	0.000
	Max	27.60	55.93	60.00	53.20	45.05	31.99	16.73	60.00
	Weighted mean	9.15	14.64	20.47	12.68	9.97	13.99	8.17	13.79
	Weighted CV	0.68	0.79	0.93	0.81	1.13	0.86	0.85	0.87

Difference		-5.5%	-1.8%	-20.7%	2.6%	-10.9%	-0.7%	-18.0%	-3.6%

For the thinner areas, generally at the longitudinal extremities of the zone, CCAF mining will be utilized. An example of cut and fill stoping is shown in Figure 16-4 below. Although more expensive than LH mining, it is more selective and produces less dilution. CCAF is chosen over mechanized cut-and-fill mining (MCF) even though it is less productive. This is because MCF would have required much more waste development to access each vertical cut of the mineralized zone. CCAF stopes are developed in 50 metre intervals from the extraction level of one stoping panel up to the next one. Stopes are up to 125 metres long and are accessed by a manway/muck raise carried from the extraction level up to the stope through the backfill. A ventilation/service raise is established in each stope up to the upper extraction level.

Equipment Type	Units	HP per Unit	HP (Total)	Availability (%)	Utilization (%)	HP (Utilized)	Air Volume (m3/sec)	Air Volume (ft3/min)
Jumbo 2 Boom	2	147	294	85	25	62	2.8	5,926
LH Drill	1	99	99	85	25	21	0.9	1,996
Mechanized Bolter	1	99	99	85	25	21	0.9	1,996
20 t Truck	3	300	900	85	90	689	30.8	65,308
3.0 m³ LHD	3	201	603	85	70	359	16.1	34,033
2.0 m³ LHD	1	70	70	85	50	30	1.3	2,822
Scissor Lift	1	147	147	90	30	40	1.8	3,765
ANFO Loader	1	147	147	90	30	40	1.8	3,765
Fuel/Lube Truck	1	147	147	90	50	66	3.0	6,275
Utility Truck	1	147	147	90	40	53	2.4	5,020
SV/Mechanic Vehicles	4	127	508	90	30	137	6.1	13,010
Subtotal:							67.9	143,913
Losses at 15%:							10.2	21,587
Total:							78.1	165,500

Description	Unit	Value
Annual Mill Feed Throughput	tpa	273,750
Operations
Crusher Availability	%	75
Grinding and Flotation Availability	%	80
Plant Daily Throughput	tpd	750
Plant Hourly Capacity	tph	39
Average ROM Feed Au Grade	g/t	10.3
Crushing
Primary Crusher	type	Jaw
Secondary Crusher	type	Cone
Tertiary Crusher	type	Cone
Fine Ore Stockpile Live Capacity	t	750
Grinding
Bond Work Index	kWh/t	19.6
Grinding Feed Particle Size F80	mm	12
Grinding Product Particle Size P80	µm	75
Mill Type		Ball Mill
Number of Mills		1
Rougher/Scavenger Flotation
Residence Time	mins	37.5
Number of Cells		6
Cell Type		Tank
Cell Volume	m3	12
Rougher/Scavenger Mass Pull	%	26.5
Cleaner Flotation
Residence Time	mins	27.5
Number of Cells		2
Cell Type		Tank
Cell Volume	m3	7.6
Final Concentrate Mass Pull	% of mill feed	15.2
BIOX™ Leaching
Plant Capacity	tpd	105
Residence Time	days	6
Number of Tanks		6
CIL and Carbon Desorption
CIL Residence Time	hours	24
Number of CIL tanks		6
Carbon Concentration	g/L	20
Estimated Carbon Loading	g/t	950
Elution Strip Rate	strips per week	7
Cyanide Destruction
Method		ASTER™ and SO2/Air
Reagents Used		Molasses, Na2S2O5, NH4H2PO4
Tailings Thickener
Target U/F Density	%	60
Number of Thickeners		1
Tailings Filtration
Filter Type		Ceramic Vacuum Disc
Design Rate	t/h/m2	0.5
Target Moisture Content	%	16
Availability	%	85
Number of Filters		2

Process	Reagent/Consumable	Consumption (kg/t Mill Feed)
Crushing	Jaw Crusher Liner	0.020
Crushing	Cone Crusher Liner	0.016
Grinding	Ball Mill Liners	0.109
	Grinding Media (Balls)	1.000
	Re-grind Media (Balls)	0.015
Flotation	Sodium Sulphide (Na2S)	0.250
	Potassium Amyl Xanthate (PAX)	0.101
	MIBC frother	0.028
	Copper Sulphate (CuSO4)	0.250
	Sodium Silicate (Na2SiO3)	0.013
BIOX™	Limestone (CaCO3)	46.232
	Lime (Ca(OH)2)	6.863
	Nutrients	1.782
	Antiscalant	0.004
	Corrosion Inhibitor	0.004
	Biocide	0.004
	Defoamer	0.001
	Flocculant	0.004
CIL/Elution	Sodium Cyanide (NaCN)	1.5
	Activated Carbon	0.030
	Defoamer	0.038
	Lime (Anhydrous) (CaO)	4.0
	Hydrochloric Acid (HCl)	0.076
	Sodium Hydroxide (NaOH)	0.198
Detox/ASTER™	Ammonium Hydrogen Phosphate (NH4H2PO4)	0.023
	Molasses	0.015
	Sodium Metabisulphite (Na2S2O5)	1.0

Fish and fish habitat have been characterized in the project area several times beginning in 1997. Spawning and rearing habitat that supports both sea-run and freshwater salmonid species including chinook, coho, pink, and sockeye salmon. The mainstem Tulsequah River has been shown to afford poor fisheries values due to low water temperatures and lack of in-stream cover typical of glacial runoff systems. Based on observations, Whitewater Creek in the project area contains fish rearing and spawning habitat superior to the Tulsequah River and therefore remains critical habitat necessary for sustaining populations (GLL 1997). Knowing this, for a project to be deemed acceptable to government and other stakeholders, future development will have to carefully consider the streams and the riparian habitat which encompasses them.

Canarc retained URS Canada Inc. to assess the acid rock drainage and metal leaching potential of major rock units anticipated to be exposed. Studies were completed by static testing conducted on 27 “fresh” rock drill core samples, collected from sections of the C vein, and flotation testing was conducted on 5 test tailings samples. A criterion of 2.0 weight % sulphide Sulphur has been developed to distinguish between non-acid generating and potentially acid generating materials. For waste rock, sampling indicated, generally, that the rock is non-acid generating in nature, due to high carbonate content. It should be noted there were some localized areas that showed some potential for acid generation. For tailings, based on static testing of 5 samples, tailings are not expected to be acid generating. This was indicated by sulphide Sulphur content being typically less than 0.3% and the corresponding carbonate content being greater than 3.5%.

The New Polaris Project lies within the area of the Atlin/Taku Land Use Plan (2011). The plan allows for Resource Development, including mining, contingent upon meeting specific requirements of Management Plans for specifies issues, including those of particular interest to the local First Nations. There are 7 objectives set in the affected land use plan concerning mineral extraction/exploration with respect to engaging with stakeholders. Above all, but not to reduce significance from others, Objective #2 states that a proponent must ensure mineral exploration and development are undertaken in a socially and environmentally responsible manner. It will be extremely important to respectfully engage with local First Nations and community stake holders to establish a positive relationship and to understand concerns with the intention to incorporate them into design, operations and closure.

Table 21-13	Build-Up of Camp Operating Costs	175
Table 21-14	Power Generation Cost Estimates	177

Table 22-1	Inputs for Economic Analysis	178
Table 22-2	Summary of Economic Analysis	179

Table 26-1	Recommended Programs for Next Level of Study	189

Confining Shape Target Grade (g/t Au)	In Situ Tonnage (tonnes)	In Situ Au Grade (g/t)	In Situ Au Content (Oz.)
2.0	1,880,000	10.0	605,000
3.0	1,798,000	10.4	599,000
4.0	1,687,000	10.8	586,000
5.0	1,556,000	11.3	567,000
6.0	1,403,000	12.0	540,000
7.0	1,260,000	12.6	509,000
8.0	1,105,000	13.3	472,000
9.0	947,000	14.1	428,000

Confining Shape Target Grade (g/t Au)	In Situ Tonnage (tonnes)	In Situ Au Grade (g/t)	In Situ Au Content (Oz.)
2.0	1,639,000	9.5	502,000
3.0	1,582,000	9.8	497,000
4.0	1,483,000	10.2	485,000
5.0	1,351,000	10.7	464,000
6.0	1,223,000	11.2	441,000
7.0	942,000	12.5	380,000
8.0	753,000	13.8	334,000
9.0	653,000	14.6	306,000

Area	Capital Estimate (M$)
Overall Site	$12.8
Mining	$20.0
Plant	$39.8
Waste/Tailings	$6.8
Site Services	$4.3
Surface Mobile Fleet	$2.5
Buildings and Facilities	$8.4
Total Direct Costs	$94.5
Total Indirect Costs	$25.2
Contingency (20%)	$23.9
Total	$143.7

Area	Operating Estimate	Units
Mining	$65.25	$/t milled
Processing	$62.70	$/t milled
G&A	$37.00	$/t milled
Total	$164.95	$/t milled

Mill Feed Production	2,306,000 tonnes (subset of Mineral Resource)
Production Rate	750 tonnes per day
Production Grade	10.3 grams per tonne
Recoveries	90.5% gold into doré
Average Output	80,000 oz gold per year
Mine life	8.7 years

The Base Case Financial Parameters:
Gold Price	US$ 1300 per oz
Exchange Rate	US$ 1.00 = C$ 1.30
Capital Cost	$143.7 million
Cash Cost All -in-Sustaining Costs	US$ 433 per oz US$ 510 per oz

Cash Flow (LOM)	Pre-Tax After Tax $554.0 million $414.4 million
NPV (5%)	$384.8 million $280.4 million

Internal Rate of Return	Pre-Tax After Tax
Internal Rate of Return	47.2% 37.7%
Payback Period	2.3 years 2.7 years

Year	Zone	No. of Holes	Metres
1988	Y vein	8	1028
1989	Y vein	19	4078
1990	C vein	10	2862
1991	Y & C veins	11	3333
1992	Y& C veins	23	6378
1993	C vein	8	1301
1994	C & Y veins, North Zone	30	5235
1995	North Zone	20	7600
1996	Underground	24	3205
1997	Underground	49	8869
1998	No drilling	0	0
1999	No drilling	0	0
2000	No Drilling	0	0
2001	No drilling	0	0
2002	No drilling	0	0
2003	C & AB veins	3	1530
2004	C vein	7	1651
2005	C vein	8	2357
2006	C vein	72	24801
Total		292	74,228 m

	MEASURED RESOURCE			INDICATED RESOURCE
Au Cutoff	Tonnes > Cutoff	Grade>Cutoff	Contained Metal	Tonnes > Cutoff	Grade>Cutoff	Contained Metal
(g/t)	(tonnes)	Au (g/t)	Au (ozs)	(tonnes)	Au (g/t)	Au (ozs)
6.00	271,000	11.89	104,000	1,017,000	12.71	416,000
7.00	233,000	12.77	96,000	910,000	13.45	393,000
8.00	203,000	13.54	88,000	806,000	14.22	368,000
9.00	173,000	14.42	80,000	696,000	15.11	338,000

	MEASURED PLUS INDICATED RESOURCE			INFERRED RESOURCE
Au Cutoff	Tonnes > Cutoff	Grade>Cutoff	Contained Metal	Tonnes > Cutoff	Grade>Cutoff	Contained Metal
(g/t)	(tonnes)	Au (g/t)	Au (ozs)	(tonnes)	Au (g/t)	Au (ozs)
6.00	1,288,000	12.54	519,000	1,628,000	12.15	636,000
7.00	1,143,000	13.31	489,000	1,473,000	12.74	603,000
8.00	1,009,000	14.08	457,000	1,340,000	13.27	571,000
9.00	870,000	14.97	419,000	1,149,000	14.06	519,000

Year	Supervisor	Drilling Contractors	Laboratory	Underground
1988	Cloutier		TSL
1989	Cloutier		TSL
1990	Cloutier		TSL
1991	Marriott	Arctic	Min En/Chemex
1992	Marriott	Arctic	Chemex
1993	Marriott/Moors	Arctic	Chemex
1994	Moors	Arctic/Falcon	Chemex
1995	Moors	Arctic/Falcon	Chemex
1996	Karelse/Watkinson	Advanced	Northern	Main Street
1997	Karelse/Watkinson	Advanced	Northern	Main Street
1998	-
1999	-
2000	-
2001	-
2002	Moors
2003	Moors	Hy-Tech	ALS Chemex
2004	Moors/Aspinall	Hy-Tech	ALS Chemex
2005	Moors/Aspinall	Hy-Tech	ALS Chemex
2006	Moors/Cote	Hy-Tech	ALS Chemex

HOLE-ID	Easting (m)	Northing (m)	Length (m)	Dip	Azimuth
03-P01	1569.72	1021.08	365.46	-70	345
03-P02	1513.33	1021.08	426.72	-70	330
03-P03	1591.06	1255.78	768.71	-60	260
04-1707E1	1706.88	873.25	259.08	-75.4	358.8
04-1707E2	1707.44	873.25	280.11	-83	359
04-1737E1	1737.63	873.16	287.43	-83	359

PT9741	1594.93	1097.74	155.14	-42	130
PT9742	1593.46	1096.88	247.65	-37	188
PT9743	1592.92	1096.85	246.58	-36	210
PT9744	1594.71	1097.4	290.17	-38	134
PT9744A	1594.71	1097.4	136.55	-35.5	141.5
PT9745	1593.1	1096.85	219.46	-26	193
PY06B	1769.36	1406.96	295.05	-60	273
PY07A	1772.11	1406.96	387.1	-67	275

P95C42	456.47	457.54	1.10	13.13
P95C43	482.04	484.51	2.47	15.89
P95C44	586.50	589.57	3.07	18.76
P95C44	640.08	641.70	1.62	12.41
P95C44	692.38	694.15	1.77	28.06
PC25A	137.31	141.00	3.69	5.60
PC44A	586.40	591.22	4.78	16.00
PC44A	726.40	730.60	4.18	6.00
PC9001	206.35	213.79	7.44	11.80
PC9002	277.31	279.96	2.65	29.30
PC9003	215.59	222.50	6.92	14.80
PC9003	227.17	232.87	5.70	31.20
PC9004	221.25	225.25	3.99	16.10
PC9005	189.43	190.96	1.52	18.90
PC9007	160.90	162.31	1.40	15.20

Parameter	Result
Population	56
Minimum value	0.05
Maximum value	0.57
Mean value	0.07
Standard Deviation	0.08
CV	1.12

Parameter	Result, First Sample	Result, Duplicate Sample	Result, Sample Difference
Population	45	45	45
Minimum value	0.025	0.025	-1.1
Maximum value	19.85	27.1	7.25
Mean value	1.52	1.81	0.30
Standard Deviation	3.72	4.76	0.18
CV	2.45	2.63	4.09

Standard	Mean (Au g/t)	Standard Deviation	Upper Range	Lower Range
PM 165	6.51	0.10	6.71	6.31
PM 415	2.37	0.12	2.49	2.13
PM 916	12.7	0.09	12.9	12.5

Parameter	Chemex Sample 1	Difference Chemex 1 vs. Acme	Acme	Difference Acme vs. Chemex 2	Chemex Sample 2	Difference Chemex 1 vs. Chemex 2
Population	30	30	30	30	30	30
Minimum value	0.0	-9.43	0.01	-0.31	0.0	-9.4
Maximum value	40.7	0.36	42.11	2.04	41.4	0.8
Mean value	7.49	-0.50	7.99	0.19	7.80	-0.31
Standard Deviation	10.47	1.82	11.31	0.43	11.08	1.76
CV	1.40	3.65	1.42	2.25	1.42	5.70

Method	Au	Ag	As	Sb	SiO2	Fe2O3	S	MgO	Al2O3	CaO	Cu	Zn
Method	g/mt	g/mt	ppm	ppm	%	%	%	%	%	%	ppm	ppm
FA/AAS	19.48	< 1.71
ICP			28,350	127
XRF			24,500	300	45.1	9.8	1.92	9.79	13.7	14.1	169	63

Mineral	Approximate Weight (%)
Dolomite	32
Mica/Illite	32
Quartz	23
Arsenopyrite	5
Pyrite	< 5
K-Feldspar	< 3
Unidentified	< 5

Method	Au	Ag	As	Sb	ST	S-2	SSO4	CORG	CT
Method	g/mt	g/mt	%	%	%	%	%	%	%
FA/AAS	9.66	1.0
ICP			0.45	0.78	3.18	3.18	< 0.01	0.33	3.7

Composite #	# of Samples	Total Weight (kg)	Average Au Grade (g/t)	% S Total	% As Total
1	116	304	9.75	2.42	1.62
2	84	150	20.44	2.26	1.74
3	82	97.5	13.92	2.53	1.74
Combined	282	551.4	13.4	2.40	1.67

Test #	Comments	Recovery (% of initial feed)			Assay Value
Test #	Comments	Wt.	Au	As	Au (g/t)	As (%)	CO2 (%)
1	PAX/MIBC	20.7	83.6	83.5	76.63	9.10
2	Add Na2S, CuSO4	18.1	93.8	93.9	94.24	11.88
3	Lower pH	21.0	87.0	86.1	78.08	10.05
4	Rougher Conc. Gen.	21.0	90.1	89.1	81.96	10.85	7.47

Sample Leached	Calc. Head Grade (g/t)	Au Extraction (%)	NaCN Consumption (kg/t)
Test 4 Rougher Tails	2.13	23.3	3.88
Test 7 Rougher Tails	0.92	29.5	4.139
Test 8 Cleaner Tails	13.15	17.9	4.696

Lab Assayed	Recovery (% of initial feed)					Assay Value
Lab Assayed	Wt.	Au	As	S	CO2	Au (g/t)	As (%)	S (%)	CO2 (%)
RDI	19.5	98.6	88.1	98.1		91.98	12.80	21.1
Mintek						89.9	14.3	19.7*
RDI						91.84	14.72	20.3	3.4
Oxidor						93.96

Test	Dissolution (%)		Au Extraction (%)	NaCN consumption (kg/t)
	Sulphide	Arsenic
Direct CIL	0	0	17.6	2.7
Dilute HNO3 + CIL	26	42.5	60.3	4.85
Conc. HNO3 + CIL	96.3	97.6	71.9	4.9
			93.7	49.3

	Baseline	180°C POX			200°C POX
Oxidation Time (mins)	0	30	60	90	30	60	90
H2SO4 consumption (pre-treatment, kg/t)		145	136	138	138	121	125
% Sulphide oxidation		98.8	99.4	100	95.1	100	100
% gold recovery(48-hour)	9.4	94.7	98.4	98.2	83.7	98.7	98.0
NaCN consumption (48-hour, kg/t)	1.33	1.18	1.41	1.03	0.79	1.24	1.19

Test #	P80 Size (lm)	Calc. Au, Head (g/t)	Pan Conc.
Test #	P80 Size (lm)	Calc. Au, Head (g/t)	Au Grade (g/t)	Au Recovery (%)
GF1	69	10.3	183.1	3.6
GF2	68	10.7	245.6	5.3
GF3	30	10.8	264.9	5.5

Test #	Test Conditions	Conc./Tails	Mass Pull (%)	Recovery (%)				Grade (g/t, %)
Test #	Test Conditions	Conc./Tails	Mass Pull (%)	Au	As	Sb	ST	Au (g/t)	As (%)	Sb (%)	ST (%)
F9LC	Rougher: 250 g/t Na2S, 50 g/t PAX, 12 g/t MIBC Scavenger (x2): 50 g/t PAX, 12 g/t MIBC, 250 g/t CuSO4 Cleaner (x3): 500 g/t Na2SiO3, 15 g/t PAX, 27 g/t MIBC	Final Conc.	12.8	91.8	90.8	86.9	94.0	72.3	10.4	5.34	22.7
		Cleaner Tails	12.0	2.6	2.9	7.8	2.1	2.14	0.36	0.51	0.53
		Bulk Tails	75.2	5.7	6.3	5.3	3.9	0.76	0.12	0.05	0.16
		Head	100	100	100	100	100

BAT #	BIOX™ time (days)	Sulphide Oxidation (%)	Arsenic Sol’n (%)	Iron Sol’n (%)	Au Dissolution (%)		Consumption (kg/t)
BAT #	BIOX™ time (days)	Sulphide Oxidation (%)	Arsenic Sol’n (%)	Iron Sol’n (%)	Residue	Calculated Head	NaCN	Lime
0	0	0	0	0	8.1	12.7	6.5	1.1
6	12	89.6	65.7	70.6	86.2	86.7	14.0	21.8
5	13	94.7	70.5	79.2	90.8	91.3	11.8	25.8
2	14	93.9	72.9	73	90.2	89.3	13.8	30.7
7	17	97.9	73.9	77.6	90.0	89.2	11.7	21.4
4	20	97.3	75.2	74.9	93.7	93.3	14.2	28.1
1	22	99.0	69.0	74.3	96.3	96.0	11.3	22.1
3	22	99.1	66.4	73.3	95.1	94.9	13.2	24.0
Final	22	99.05	67.4	73.8	95.7	95.45	12.25	23.05

		Neutralization Products
Test #	Test Conditions	Solids (%)		Solution (mg/L)		TCLP (mg/L)
Test #	Test Conditions	Fe	As	Fe	As	Fe	As
1	Lime + Limestone slurry	15.7	5.8	0.1	0.3	< 0.1	10
2	Lime slurry	14.8	6.1	0.4	0.4	< 0.1	3.9
3	Lime + Limestone + Ferric Sulphate	16.1	5.0	< 0.1	0.4	< 0.1	2.3

	Flocculant Tested (100 g/t)
	Magna 10	Magna 333	Magna 345	Magna 455	Magna 336	Magna 156	Magna 1011	Magna 405
Settling time (sec)	30	30	30	30	30	30	30	30
Displacement (mm)	5	3	3	3	3	3	5	13
Settling Rate (m/h)	0.60	0.36	0.36	0.36	0.36	0.36	0.60	1.56
Overflow Clarity	Cloudy	Cloudy	Cloudy	Cloudy	Cloudy	Cloudy	Cloudy	Clear

	Flocculant Tested (100 g/t on 7.6% solids)
	Magna 10	Magna 333	Magna 345	Magna 455	Magna 336	Magna 405
Settling time (sec)	45	45	45	45	45	45
Displacement (mm)	7	6	14	10	7	6
Settling Rate (m/h)	0.56	0.48	1.12	0.80	0.56	0.48
Overflow Clarity	Clear	Clear	Clear	Clear	Clear	Clear

Area	Recovery (%)
Sulphide Flotation	94.9
BIOX and CIL Leach	95.6
Carbon Loss	0.1
EW	99.9

Indicated
Confining Shape Target Grade - (g/t Au)	In Situ	In Situ Grades
	Tonnage	AU	Au
	(Ktonnes)	(g/t)	(koz.)
2.0	1,880	10.0	605
3.0	1,798	10.4	599
4.0	1,687	10.8	586
5.0	1,556	11.3	567
6.0	1,403	12.0	540
7.0	1,260	12.6	509
8.0	1,105	13.3	472
9.0	947	14.1	428
Inferred
Confining Shape Target Grade - (g/t Au)	In Situ	In Situ Grades
	Tonnage	AU	Au
	(Ktonnes)	(g/t)	(koz.)
2.0	1,639	9.5	502
3.0	1,582	9.8	497
4.0	1,483	10.2	485
5.0	1,351	10.7	464
6.0	1,223	11.2	441
7.0	942	12.5	380
8.0	753	13.8	334
9.0	653	14.6	306

Source	Drillholes	Total Length (m)	Au Assay Intervals above Zero	Au Assay Length (m)	Percent Assayed and above Zero
Project Total	1056	106,458.6	11,590	11,850.7	11%
Block Model Total	709	85,546.0	9,645	8,481	10%
Within block model and Since 1988	286	73,898.1	7,895	8,393.2	11%
Domain Total	173	50,704.0	1,357	1,101.7	2%
Domain Total since 1988	143	47,042.4	1,287	1,033.2	2%